Apparatus for the continuous smelting and converting of copper concentrates to metallic copper



United States Patent Inventors Nickolas J. Themelis [56] References Cited Bumsflelt Quebec; uwmzo STATES PATENTS I N 535%" Mmmammbectcanada 2,034,071 3/1936 Wickland 266/11x P 2,438,911 4/1948 Gronningsaeter.... 266/11x SW13 3 326 671 6/1967 Worner 7s/73x Division of Ser. No. 423,257, now Patent No. 3,437,475, dated Apr. 8, 1969 Primary Examiner-J. Spencer Overholser p d N 24, 1970 Assistant Exam iner-J ohn E. Roethel Assignee Noranda Mines, Limited Attorney-(3118mm Darby & Cushman Toronto, Ontario, Canada Pnomy :gtfii 1964 ABSTRACT: Apparatus is provided for a continuous process 917 049 for smelting and converting copper concentrates to metallic copper and cleaning the slag resulting therefrom. The apparatus comprises a generally horizontally disposed furnace including a smelting zone, a converting zone and a slag-settling zone; means for rotating the furnace about its longituggg g dinal axis; a charging port formed in said furnace; means for To METALLIC COPPER ATES introducing an oxidizing gas into the furnace; and wherein the means is shaped and dimensioned to ensure the introduction 6 Clams 2 Drawing of a sufficient volume of oxidizing gas to effect gradual oxida- U.S.Cl 266/36, tion of the matte and subsequent conversion to metallic 263/33 copper; a copper-settling section provided in the base area of Int. Cl.....'. C2lc 5/50: the furnace; an area of the furnace which is remote from the F27b 7/06 charging port being'shaped to form a reducing section for slag Field ol'Search 75/73, 30, resulting fr m h pr n i rg p r f r th opp r 34,40,5l,74, 75;266/ll,35,36R, 18,34; 263/33 and slag.

a dc Flt/X smcz m 4 5 1-1 Co/vce/vrznrs Patented Nov. 24, 1970 YQQQL APPARATUS FOR THE CONTINUOUS SMELTING AND CONVERTING OF COPPER CONCENTRATES TO METALLIC COPPER This is a division of U.S. Pat. application Ser. No. 423,257, filed Jan. 4, 1965, now U.S. Pat. No. 3,437,475.

This invention relates broadly to the smelting and conversion of copper concentrates to metallic copper. More particularly, the invention is directed to a process and apparatus for the continuous smelting and conversion of copper concentrates to metallic copper.

The invention described herein is intended to replace the conventional batch-blowing smelting process (which will be described hereafter for purposes of comparison) by a continuous copper reduction process such that the smelting and converting stages occur partly simultaneously and partly in sequence in a single reactor vessel in which the concentrates are introduced continuously at one end, while the slag and copper metalare removed continuously at the other end of the vessel.

The process ofthe present invention also includes thetreatment of slag by reducing gases or other means with the purpose of decreasing the copper content to an accepta le level, either in an appropriate extension of the reduction r actor or in a separate vessel adjacent to the reactor, such that'the slag can flow by gravity from the reactor vessel to the settling vessel.

In addition the invention embodies the treatment of the produced copper to anode copper on stream" first by oxidizing and then by reducing gases either in an appropriate extension of the reactor proper or in one or more holding furnaces :heated by induction or other means.

The invention therefore comprises the continuous smelting converting and reducing of copper concentrates and subsequent treatment of the copper and slag streams which, due to the continuity of 'the process, occupy a small volume er unit time and may therefore be treated on stream to produce a relatively copper-clean slag stream and a copper stream suitable for anode casting.

The invention also utilizes certain aspects of present-day copper technology, in conjunction with new concepts of on stream gas-liquidprocessing which are up to the present time used mostly in chemical processing. The invention also includes the concept of removing copper from molten silicates by subjecting the slags to contact with highly reducing atmospheres and allowing the reduced copper to settle out of the slag.

It is known that for many years considerable thought and effort has been directed to the formulation of means for continuously converting copper concentrates but so far no satisfactory or commercially workable process has been developed. By way of explanation and comparison itmay be noted that'the conventional noncontinuous or batch copper smelting and converting processinvolves meltingof the concentratesand flux in a reverberatory or blast furnace wherein two separate layers are formed-a heavier one of matte '(Cu S-FeS') and supernatant layer of slag. The supernatant layer is allowed to settle and is cleaned of most of its copper content. The matte from the reverberatory furnace is then conveyed to-the convertervessel where it is subjected to two stage air-oxidationTeaction. In the first stage of the converting reaction, oxygen reacts withFeS as follows:

FeS+ 1%O :FeO+ S (1) Any Cu S which may be oxidized to Cu O reacts immediately with FeSaccording to:

CLlgO-FFGSZ-CllgS-i-FGO (2) Silica flux is addedto'the converter continuously to form ironsilicateslag with theFeO produced by reactions (l) and 2Fo0+ SiO FJFCOJSiO (3) The slag produced in the first stage of air-blow is then skimmed from the converter and transferred to the reverberatory furnace where mixing and interaction with the furnace bath lowers its copper content from 23 percent Cu to about 0.35 percent in the reverberatory slag.

The Cu S (white metal) which has remained in the converter is then subjected to a second blow which is believed to result in the following reactions:

with the net result of producing metallic copper.

There are technical problems which have prevented the earlier development of a continuous converting process. For example, the thermodynamics of copper conversion require that the process be carried out in two steps or stages. The first stage involves oxidation of FeS and the second oxidation of Cu S to Cu. It had previously been concluded that the batch operation of the converter lends itself best to these reactions. In addition, it had been found that during the second stage of the process involving conversion of Cu S to Cu what may be termed the end point of the conversion was frequently overstepped and a certain amount of copper oxidized as Cu O. This Cu O dissolves in any slag which may be present in the converter at the time.

Therefore, it had been considered that in order to keep the amount of copper in converter slags relatively low it was necessary to have very little or no slag during the second blow. This is a requirement which has been avoided by the present continuous process where the slag from the first stage or blow is present during the second blow.

The need for an efficiently operating process and apparatus for continuously smelting and converting copper concentrates to metallic copper is therefore apparent. The broad object of the invention is to fill this need.

According to the invention a process and apparatus for the continuous smelting and converting of copper concentrates to metallic copper in a single furnace involving a gradual and sequential reaction along the length of said furnace includes the steps of (a) feeding flux and the concentrates to be converted into said furnace; (b) smelting the concentrates and flux; (c) controlling the resultant flow of matte and slag in the furnace as it flows towards tapping ports formed therein; (d) introducing an oxidizing gas into the matte sufficient to effect a gradual oxidation of the ferrous sulphide; (e) continuing to introduce said oxidizing gas into the resultant white metal in a volume sufficient to effect the gradual oxidation of the copper sulphide to metallic copper; (f) allowing the copper to settle and thereafter (g) drawing off metallic copper.

FIG. 1 is a side view of a generally horizontally disposed reaction vessel of the present invention. 1

FIG. 2 is an end view of the vessel of FIG. 1.

In FIG. 1 of the drawings the vessel designated generally at A has an outer casing 1. A charging port 2 is formed in the casing 1 for the introduction of the concentrates to be reduced. A burner 3, whichmay be conventional in structure and operation, is also located at the charging end of the vessel. An exhaust stack 4 is provided in the normally upper portion of the vessel. It will be noted that the horizontal length of the vessel is divided up into distinct reaction zones. Firstly, there is the zone for concentrate smelting and matte blowing X, secondly a zone Y where white metal blowing takes place, thirdly the copper settling zone Z and finally the slag reduction zone R and settling zone S atthe discharge end of the vessel. It should be noted, however, that the separation of these zones is only broadly indicated. They are not separate and distinct reaction zones in the sense of being physically divided by partitions or the like. Tuyeres 5 are located along the length of the vessel as illustrated. These tuyeres provide for the continuous introduc tion of air to ensure that the continuous aspect of the process is accomplished. Inlet means for a reducing gas is shown at FIG. 2 at 6. The additional heat required in the slag settling zone may be provided by means of either burner 7 or by supplying a sufficient amount of air through inlet 8 to burn off the excess reducing gas emerging from the slag layer. The furnace may be rotated about its longitudinal axis by any conventional means, a suitable means may consist of riding rings a resting on rollers b and a ring gear c driven by a gear d connected through a speed reducer e to a motor f.

Although the present process is continuous, a gradual and sequential reaction is maintained from the first to the second stage of blowing. This is achieved by controlled flow and oxidation of the liquid matte stream as it moves from the feeding end of the furnace towards the tapping end thereof.

Thus, the transition from the first stage involving FeS oxidation to the second stage involving Cu S oxidation is not sudden-as it is in the presently used batch operation processesbut rather -it is a gradual transition; the FeS is slowly oxidized out of the matte until it reaches the second section of the furnace where only Cu S is present.

The invention is not limited to any particular or critical shape or size of reactor vessel or equipment and can also be applied to other types of liquid-gas reactions at high temperatures such as certain other sulphides. The following specific example is given of copper production and the continuous method of the present invention and should be considered in the light of the drawings.

PRODUCTION OF COPPER Basis: 200 tons of copper produced/day Material balance:

Wet concentrates-32.1 tons/hr.

Assumed composition.26% Cu; 30% Fe;

33% S; 4% SiOz; 7% H O. Tons of matte produced (29% Cu)28.6 tons/hr. Volume of matte (S.G.=4.1)225 cu. ft./hr. Flux addition-7.5 tons SiO /hr. Tons of slag produced: 22.6 tons/hr.

Composition45% Fe (58% FeO); 30% SiO 12% A1 ZnO, etc. Volume of slag (s.g.=3.6)=202 cu. ft./hr. Tons of White metal 0112s produced-10.45 tons/ hr.

Volume (e.g.=5.9)56.8 cu. ft./hr Tons of metallic copper produced8.35 Volume (s.g.=8)33.5 cu. ft./hr. Air requirements:

(a) First stage blow:

FeS to FeO at 95% oxygen utilization 15,500 c.f.m. air.

tons/ hr.

Volume of air per 2-inch tuyere-250 c.f.m. of air.

(Present operation praetice500 c.f.m.)

Number of tuyeres required-62.

Length of reactor required (at 6-inch centres for tuyere as present operating practice)32 ft.

(b) Second stage blow:

Cu S to Cu at 95% oxygen utilization 4,000 c.f.m. of air.

Number of tuyeres-16.

Length of reactor required (6-inch centres)9 ft.

Residence Times of Materials in Reactor The reactor vessel A may be very similar in design to the conventional Peirce-Smith converter. Therefore, full conversion will be achieved in the same length of time as obtained in present batch operation. Consequently, the residence time of material is considered to be about 2 hours in the first-stage flow zone X and approximately 0.5 hours in the second-stage blow zone Y. in addition, a residence time of about 1 hour is undertaken for the settling of copper in zone Z and about 0.5

hours for the reducing blow in the slag and 1 hour for the settling of the slag in zone S. The latter figure has been based on experiments on the effect of settling on copper slags. As indicated there are no separate blows as such but rather separate reactions or stages in the process resulting from the continuous blowing and introduction of air through the tuyeres 5.

5 practice.

Heat Balance (Per hour of operation):

(a) Matte-blowing zone: B.t.u./hr.

Heat content of liquid matte at 2,200 F 22, 500, 000 Heat content of liquid slag at 2,200 F. (including 10 tons/hr. of flux) s 29, 400, 000 Heat content of air input at 2,200 F s 40, 900, 000 Heat content of water vapour at 2,200 F 9, 700, 000 Heat losses (operating experience, 30 ft. converter) 6, 000, 000

Total Heat required 107, 600, 000

Heat available from FeS reaction and formation of silicate slag 76, 500, 000 Auxiliary heat supply at feeding end of reactor (e.g. 3 tons of coal at 42 available heat) 31, 100, 000

Total available heat 107, 600, 000

(b) White metal-blowing zone:

Heat content of added silica flux (2.5 tons) at 2,200 F 3, 000, 000 Heat content of air at 2,200

10, 000, 000 Heat losses s- 1, 500, 000

Total heat required s 14, 500, 000

Heat of Cu S reaction with oxygen (860 B.t.u./lb. Cu). 14, 500, 000 Therefore, total available heat 14, 500, 000

(c) Copper-settling, slag-reducing,

and slag-settling zones (estimated total length required: 30 ft.) Heat content of reducing gas (55 c.f.m. of natural gas) 500, 000 Heat losses 6, 000, 000

Total heat required 6, 500, 000

Auxiliary heat source (burner on tapping end of reactor, 230 c.f.m. of natural gas at 50% available heat) 6, 500, 000

Total fuel to process:

3 tons of coalX 24,000,000 72, 000, 000 285 c.f.m. 60X 1,000 17, 100, 000

A description of process may now be considered with reference to specific example earlier discussed.

The wet (or dry) concentrates'are premixed with the required proportion of the silica flux and fed into the front end of the bath of the reactor vessel A by means of any suitable feeding mechanism such as a vibrating screw feeder, a belt feeder, a Garr gun or the like. Experiments have shown that when the concentrates reach a temperature of about 1,000- llO'F., reaction with the airstream from tuyeres com, mences and the exothermic heat of reaction brings the concentrates to the smelting andj converting temperatures.

The heat balance has shown that most of the required heat in the reactor is supplied by the exothermic heat of the converting reactions. An additional heat supply in the form of a gas, oil, or coal burner may be located at the feeding or charging end of the reactor.

The reactor vessel A comprises a cylindrical vessel of about 1 l feet in internal diameter and 60 feet length and may be ofa construction similar to a'Peirce-Smith converter with the exception of the raised hearth at the tapping end which allows the separation of the slag and white metal layers. Tuyeres 5 continuously discharging 250 c.f.m. each are located at 6-inch centres along zones X and Y of the reactor length. As in the conventional Peirce-Smith design, the reaction vessel can be rotated up to 90 so as'to expose the tuyeres, in the case ofinterruption of airflow. It must be visualized that raw materials are fed continuously and molten slag and copper are withdrawn continuously from the reactor furnace. Consequently, materials flow by gravity slowly past any particular point in the reactor (from left to right on FIG. 1 of the drawing).

As the concentrates melt, two layers are formed, one of slag and the other of matte, just as in the conventional reverberatory furnace. However, as the matte layer flows toward the discharge end 7 of the vessel A, it is continuously subjected to reaction with an airstream injected through the tuyeres, by continuous introduction of air. In practice a minimum of about 10 psi. will be used. It is not intended to suggest that absolutecontinuity of an introduction is essential. Some small stoppage in air introduction is not critical but the continuous aspect of an introduction is to be distinguished from the separate air blows used in the batch process of the prior'art. This interaction primarily oxidizes the ferrous sulphide content of the matte to ferrous oxide which combines with the excess silica in the slag to form a silicate slag. As noted earlier, silica is introduced in the furnace in admixture with the concentrates. An additional amount isadded by suitable feeders near the central area of the furnace.

The converting reaction described above, corresponds to the first-stage blow in conventional copper converters with the exception that in the present invention a layer of slag exists I above the matte at all times. In thisrespect, the reaction envisaged is identical to thefinal minutes of a first-stage blow when a certain layer of slag is also found on top of the white metal-matte layer.

As a consequence of this reaction between matte and airstream, the former becomes continuously depleted of FeS until at some downstream point, the matte has been converted to white metal (Cu S).

White metal is then subjected to reaction with air through the tuyeres 5 and the equivalent of the second-stage blow" of conventional converter operation starts taking place. This part of the reaction is identical to the second-stage blow", with the exception that a layer of slag exists above the white metal.

In conventional practice, it is highly undesirable to have slag in equilibrium with white metal during the blow, since a relatively large amount ofcopper finds its way into the slag. However, this aspect of previous practice need not be followed in the present case since at a later stage in the reactor the slag is subjected to a copper-cleaning treatment. Thus, it is of no special consequence if the slag contains for example 1.5 percent Cu or 5-.-6 percent Cu. Thermodynamic studies have shown that even at equilibrium conditions between metallic copper and highly oxidized slag, the concentration of copper dissolved in the slag does not exceed 6-9 percent Cu.

As the white metal is subjected to interaction with air. it separates into a sulphur-rich and a copper-rich (98-99 percent Cu) phase which has a higher density and therefore precipitates as a third liquid phase toward the copper-settling section Z of the reactor vessel A.

The zone Z of the reactor vessel. A is allotted for settling of this copper which is then tapped continuously into one or more small induction furnaces where it can be desulphurized and then deoxidized on-stream prior to casting into anodes. The slag layer flows past the copper-smelting zone into the slag-treating zoneR where it is subjected to contact with reducing gases as indicated on FIG. I, which gases are introduced through a series of tuyeres 6. A final slag-settling zone S isprovided after the reducing blow, following which the slag is skimmed continuously from the furnace, and sluiced with water. Copper which settles out of the slag flows back into the copper-settling zone Z. As an alternative a separate holding furnace may be provided whereby high copper slag is skimmed from the copper settling zone Z and is reduced by subjecting it to a blow with reducing gases and allowed to settle in order to recover its copper content in the form of a settled high grade matte tapped from the bottom of such holding furnace.

Laboratory and pilot plant experiments conducted at Noranda Research Centre have verified the essential parts of the invention as follows:

a. smelting-Converting Copper concentrates were mixed with sufficient silica flux to produce a fayalite (about 30 percent SiO slag on conversion. Experiments were conducted in large crucibles and also in a pilot furnace having hearth dimensions of 3 feet wide X 7.5 feet long. In both types of experiments, the blow was interrupted when the three phases (slag, white metal, copper) were found to coexist in the furnace. It is significant to note that white metal and copper were found at the settling end of the furnace, while the ferrous sulphide content of the matte had not been eliminated completely in the charge end of the furnace. Thus, a concentration gradient (1 FeS existed in the Cu S layer according tothe requirements of the process.

These tests showed that the separation of the three phases was easily achieved and that the converted copper settles very rapidly out of the white metal, forming a very distinct interface with the latter. The metallic copper contained from 98.6 percent to 99 percent copper.

b. Cleaning of Slag Tests were conducted in large crucibles by blowing the slag with reducing gas (partially oxidized methane) and allowing the copper to settle out of the slag. Experiments have shown that slags containing 25-30 percent Si0 and up to 5 percent Cu were cleaned by subjecting them to a blow with carbon monoxide to 0.6 percent Cu after treatment of only 15 minutes. Slag could be cleaned to approximately 0.35 percent after a residence time of '1 hour.

The reducing gases emerging from the slag layer may either be exhausted or burned with air introduced at the slag-reducing zone, to provide an additional source of heat in the settling end of the reactor.

In summary it can be stated that this concept of gradual converting along the axis of a horizontal reaction vessel is new to the copper-smelting field, although it has been utilized in thechemical engineering processes.

For the first time, the various forces affecting movement of slags and mattes (inertial, gravitational, buoyancy, etc.) have been studied and correlated by dimensional analysis in order to predict and control the behaviour of matte and slag in the bath.

The invention therefore presents a significant change in copper-converting which provides for the continuous converting of copper although it effectively separates the first and second blows, by progressive oxidation of the matte stream. In addition, continuous converting produces a steady stream of slag which can be treated in order to decrease its copper content. In the present invention the slag is subjected to a reducing blow and then settling which reduces its copper content to an acceptable and economic level. Thus, the slag can be then discarded, dispensing .with the need for a reverberatory furverting copper concentrates to metalliccopper and cleaning the slag resulting therefrom comprising:

a. a generally horizontally disposed furnace including a smelting zone,-a converting zone and a slag-settling zone;

b. means for rotating said furnace about its longitudinal axis;

c a single charging port formed in said furnace;

d. means for introducing an oxidizing gas into said furnace;

e. said means being shaped and dimensioned to ensure the introduction of a volume of oxidizing gas, sufficient to effect gradual oxidation of the matte and subsequent conversion to metallic copper;

f. a copper-settling section provided in the base area of said furnace;

g. an area of said furnace remote from said charging port being shaped to form a reducing section for slag resulting from said process;

. discharge ports for the resultant copper and slag;

. the special sequence of said zones being arranged from the location of the charging port in the order of smelting zone, converting zone and slag-settling zone;

j. the construction of said zones being such that the matte is permitted to flow concurrently from the smelting zone to the converting zone.

2. Apparatus as claimed in claim 1 vvherein said means for introducing an oxidizing gas into said furnace, comprise a plurality of tuyeres arranged along the length of the said furnace.

3. Apparatus as claimed in claim 1 including a separate reducing vessel for cleaning resultant slag to obtain a substantially copper-free slag.

4. Apparatus as claimed in claim 1 including the provision ofa burner at the charging end of the said furnace. said burner being adapted to introduce additional heat into the smelting zone of said furnace.

.5. Apparatus as claimed in claim 1 including means for providing additional heat in the slag settling zone of the said furnace, said means being in the form of a burner mounted at the slag discharge port of said furnace.

6. Apparatus as claimed in claim 1 including means for providing additional heat in the slag settling zone of said furnace. 

